Process for producing ductile vanadium



United States Patent 3,098,021 PROiIESS FGR PRODUCING DUCTILE VANADIUM Donald J. Hansen, Niagara Falls, and William J. West, Lewiston, N .Y., and James G. Farmer, Riverside, Conn, assignors to Union Carbide Corporation, a corporation of New York No Drawing. Filed Apr. 15, 1960, Ser. No. 22,409 4 Claims. (6!. 2l)4d4) This invention relates to a process for producing ductile vanadium.

Pure vanadium metal possesses many unique properties; these properties have been found useful in atomic energy applications. However, even small amounts of impurities such as oxygen, nitrogen or carbon severely embrittle vanadium to an extent to render cold working of the material virtually impossible. Techniques have been developed for producing a satisfactory product. However, none of these have been found wholly satisfactory from a commercial standpoint.

Vanadium may be purified by an electrolytic technique similar to that developed for titanium. This is effected by preparing a fused salt electrolyte of a lower halide of vanadium in a suitable alkali metal halide or alkaline earth metal halide, introducing a suitable cathode into the electrolyte for the deposition of vanadium thereon, and placing the vanadium metal to be purified in said electrolyte in an anodic condition relative to the cathode. Under the influence of direct current, the lower vanadium halide is reduced and deposited on the cathode while vanadium metal from the anode goes into solution in the electrolyte. A process of this type is described, for titanium, in the Bureau of Mines report by J. R. Nettle, D. H. Baker, Jr., and F. S. Wartman, entitled Electrorefining Titanium Metal, a report of Investigation 5315, February 1957.

Although a few percent of oxygen may be tolerated in the anodic vanadium material, i.e., the cell feed, amounts in excess of this, such as from inleakage of air into the electrolyte cell or by salt addition to the electro lyte bath, will contaminate the cathode product.

It is an object of this invention to provide a process for the electro-refining of vanadium metal to produce a cathode product substantially free from oxygen contamination.

Another object of the invention is to provide a process for the production, from vanadium pentoxide, of vanadium substantially free from oxygen and carbon contamination.

Other objects will be apparent from the subsequent disclosure and appended claims.

The process which satisfies the objects of the invention comprises introducing elemental aluminum into an electrolytic cell of the type previously described together with the vanadium metal to be refined. The aluminum and vanadium should be in intimate contact with each other so that they are simultaneously rendered anodic. The aluminum reacts with dissolved or combined oxygen contained in the vanadium according to the equations.

It may be seen, therefore, that the ratio of aluminum introduced into the cell to the oxygen contamination should be at least that found in aluminum sesquioxide, A1 0 if substantially all of the oxygen is to be removed. However, improvements can be obtained with lesser amounts of aluminum. Preferably a slight excess of aluminum is desired in order to accommodate removal of oxygen from the electrolyte as Well as from the anodic vanadium. However, caution should be observed that 3,098,021 Patented July 16, 1963 such excessive quantities of aluminum as will interfere with the vanadium electrorefining process are not introduced into the electrorefining cell.

While the aforementioned electro-refining process provides substantial improvement in the ductility of the vanadium cathode product, nevertheless it does not account for the introduction into the electrolytic cell of undesirable impurities in the form of reducing agent utilized in the production of the vanadium anode material from vanadium pentoxide and other oxidic materials. It has been found that the so-called aluminother-mic reaction of elemental aluminum with vanadium pentoxide to produce vanadium-aluminum may be utilized for the production of the vanadium anode source material. Reducing the vanadium pentoxide with aluminum provides an aluminum-vanadium alloy since aluminum alloys readily with vanadium. This product may be readily utilized as anode material. In the electrorefining process, any aluminum sesquioxide carried over with the aluminum-vanadium alloy, and any produced by the reaction of excess aluminum with oxygen contamination, forms an insoluble anode sludge which is readily removed from the electrolyte by standard means.

While caution was indicated in the amount of excess aluminum to be introduced into the refining cell, nevertheless the amount of aluminum which may be introduced has Wide latitude. This results from the fact that large quantities of aluminum may be alloyed with vanadium in the final cathode product without serious effect on the ductility and other desirable properties of the vanadium. Up to 10 percent aluminum may be tolerated in many cases. In general, however, the aluminum content of the final product should be maintained below about 5 percent and preferably below about 2 percent.

This is readily controlled, when utilizing the alumino thernn'c reaction for the reduction of vanadium pentoxide, by controlling the excess aluminum in the aluminothermic reaction. If the anode feed material contains an excess of aluminum greater than 7.5 percent, the aluminum is readily transferred to the final cathode product. If the amount of aluminum in the anode material is less than 0.5 percent, no substantial improvement is obtained. Excellent results are obtained when the aluminum content is about 4.5 percent to 5 percent. For an optimum balance of yield and efliciency, as well as oxygen con taminati'on reduction, a Q. percent aluminum-vanadium anode feed material is preferred.

The aluminothermic reaction is violent and highly exothermic. The slag will separate clearly from the reduced metal which forms a regulus in the lower portion of the reaction vessel. Small batches may be prepared in crucibles composed of refractory material, such :as magnesia, wherein the charge is first packed and then ignited, for example, by means of a magnesium-sulfur mixture. Large-scale reductions can be conducted in standard opentop, three-phase electric arc furnaces. The are starts the reaction but is required to add but little heat while the reacto-n is going on.

The regulus formed by the reaction is crushed preparatory to adding to the cell, in which condition it more readily dissovles by electrochemical action. Besides providing adequate surface area to accommodate adequate electric current for electrolysis, a size equivalent to approximately 2 mesh and through is convenient to handle.

The most satisfactory method of preparing the electrolyte, comprising a solution of approximately 3 percent vanadium subchloride in sodium chloride, is to chlorinate vanadium metal submerged in fused sodium chloride. As to materials, :any technical-grade sodium chloride can be used analyzing approximately 99.9 percent NaCl, balance incidental impurities. The source of vanadium may be derived from vanadium metal scrap,

such as the off-grade rejected product of the electrolysis step or the product from the aluminothermic reduction step. Using the off-grade material allows the vanadium to be recycled. Being able to recycle in this manner has obvious economic advantages. Any technical-grade chlorine gas may be used for chlorinating.

In one embodiment of the process for preparing the electrolyte, the reaction is carried out in a graphite container using an inert argon atmosphere. The desired quantity of sodium chloride is fused and held at approximately 825 C. Scrap vanadium metal in excess of that required to give the desired concentration of vanadium in the electrolyte can be introduced into the reaction chamber by placing it in a graphite container having a recessed perforated bottom, said containter being then completely immersed in the sodium chloride. Chlorine is fed through a graphite tube to the underside of the perforated bottom, the extended walls forcing the gas up through the perforations. The desired concentration of vanadium chloride in the electrolyte can be controlled by metering the total quantity of chlorine fed to the bath.

Although it is preferred to carry out the preparation of the electrolyte in a separate reaction vessel where larger, more uniform quantities can be made, the preparation can be carried out within the cell prior to the electrolysis step.

The electrolytic precedure is begun by adding a given amount of crushed anode material to the bottom of the cells, followed by a given amount of electrolyte. Care should be exercised to keep surface contamination to a minimum. In order to prevent the absorption of moisture and other atmospheric constituents, the electrolyte may be conveniently transferred from vessel to vessel by conduction in the molten state through air-tight conduits. After loading the cell, heat is applied until the temperature is raised to about 825 to 850 C. Direct current potential is applied between cathode and anode and maintained until a sufficient deposit of vanadium is built up on the cathode.

The electro-refining may be carried out in any suitable apparatus equipped for maintaining air-tight operation. It is necessary to exclude air because of the strong afiinity between oxygen and vanadium at the temperatures of operation, as well as for the aforementioned reasons. Graphite is a satisfactory material for containing the electrolyte. External or internal resistance heating of the electrolyte may be used with equal success provided means of thermally insulating, maintaining air tightness, and devices for admitting electrodes and similar accouterments are used.

Initial cathode current densities ranging from approximately 300 to 2500 amperes/square foot maintained for [from 2 to 70 hours have been applied, and current efiiciencies have been recorded ranging from 50 to 99 percent, based on :deposition [from the divalent state.

Another significant feature of the process is the provision of a means of controlling the composition of the final product. As stated previously, the aluminum and oxygen present in the anode material, derived from the aluminothermic reduction step, interact with one another in such a way that both constituents are scavenged therefrom as aluminum oxide sludge. Any excess aluminum present in the anode material will, likewise, react with any oxygen appearing in the electrolyte, either initially or as a result of leakage of air into the system. Thus, the scavenging effect of aluminum can be exploited as a means of regulating the oxygen content of the final product by making aluminum additions to the cell in the course of the operation, after first ascertaining the composition of the deposited metal by a spot check analysis. An additional advantage is that the off-grade deposited metal may be recycled to the cell to be used as new anode or to the chlorinator for feed for the electrolyte.

The following examples illustrate operation of the process of the invention:

4 EXAMPLE I Fourteen pounds of dry, chemically-pure vanadium pentoxide and 6.9 pounds of aluminum shot, representing stoichiometric amounts of the respective ingredients, were thoroughly blended. The mix was packed in a magnesia crucible and ignited using magnesium and sulfur to trigger the reaction. A regulus was formed which weighed approximately 5.7 pounds and analyzed 0.64 percent aluminum, 2.1 percent oxygen, 0.0032 percent hydrogen, and 0.14 percent nitrogen. The resulting vanadium alloy was reduced to Ar-inch pieces and finer and subsequently used as the soluble anode material in the electrolytic cell.

The electrolyte was prepared by vacuum fusing 17 pounds of sodium chloride at 850 C. in a graphite crucible. Two pounds of vanadium metal were immersed in the molten sodium chloride and chlorine gas passed through the metal to form a subvalent electrolyte. In this reaction, 1.7 pounds of vanadium metal and 2.3 pounds of C1 were consumed. Analysis indicated the electrolyte contained 8.62 weight percent lower valent vanadium. Three pounds of the above-prepared A-inch mesh-and-through vanadium metal were added to this electrolyte to function as soluble anode during the subsequent electrorefining step. The first electrolysis was started by immersing a steel cathode into the fused electrolyte and making the anode connection to the graphite container. A direct current of 49 amperes was passed for 2 hours. Vanadium metal was deposited on the cathode as dendrites, needles, and fines. After leaching in water to remove electrolyte, 68 grams of metal were obtained indicating a current efficiency of 72.3 percent, based on deposition from the valence state of V. Before beginning a second electrolysis, 67 grams of 4-inch mesh and through vanadium anode source material were added to the molten electrolyte. A current of 25 amperes was passed for 25 hours yielding 426.4 grams of metal that had deposited at a current efliciency of 72 percent. This metal contained 0.18 percent oxygen, 0.0021 percent hydrogen, and nil nitrogen and had a Rockwell A hardness of 49.7. After 453 grams of vanadium anode source material were added to the cell, the electrolysis was continued at a current of 25 amperes for 20 hours and 48 minutes. At the end of this time, 268.1 grams of vanadium were deposited at a current efficiency of 53 percent. The fourth electrolysis was started by adding 322 grams of vanadium anode source material to the cell. The electrolysis was run at a current of 9.5 amperes for 66 hours. This operation produced 483 grams of metal which had deposited at a current efficiency of 76 percent. This metal analyzed 0.21 percent oxygen, 0.0018 percent hydrogen, 0.004 percent nitrogen, and 0.007 percent aluminum and had a Rockwell A hardness of 53.

EXAMPLE II One hundred sixty grams of an aluminum-vanadium alloy, produced by the reduction of vanadium oxide with aluminum, which analyzed: vanadium84.2 percent, aluminum12.41 percent, silicon0.85 percent, iron0.9 percent, oxygen-0.14 percent, hydrogen0.00ll percent and nitrogen-0.02 percent were jaw crushed to A-inch mesh and through for use as an impure vanadium anode material. This material was added to a sodium chloridelower valent vanadium chloride electrolyte which was previously prepared by adding grams of vanadium dichloride to 520 grams of fused sodium chloride (7 percent soluble vanadium). Electrolysis was started at 7.5 amperes at 2 volts for 5% hours to give 26.4 grams of leached metal, indicating a current efficiency of 70 percent based on deposition from the divalent state. This metal analyzed: 0.08 percent oxygen, 0.01 percent nitrogen, 0.032 percent carbon, and 0.15 percent aluminum. Thirty grams of vanadium source material were added to the fused electrolyte, and the electrolysis continued at a temperature of 790 C. to 800 C. for 7 hours with a current flow of 7.5 amperes. This yielded 30.4 grams of leached metal at a current etliciency of 62 percent. The metal analyzed: 0.08 percent oxygen, 0.0l percent nitrogen, 0016 percent carbon, 0.51 percent aluminum and, after arc melting, had a hardness of Rockwell A/44 and Rockwell B/ 77 and was cold-rolled without annealing to sheet (+95 percent reduction). The electrolysis was continued after another addition of 30 grams of anode source material at current flow of 7.5 amperes at 1.4 volts for 6 hours, to give 27.5 grams of water-leached vanadium metal that had deposited with a current etfi-ciency of 63 percent. The vanadium analyzed: 0.019 percent carbon, 0.09 percent oxygen, 0.01 percent nitrogen, and 2.79 percent aluminum. The deposit was arc-melted into button form which had a Rockwell A/ 60 hardness of 49. This button was then cold-rolled to +95 percent reduc tion.

EXAMPLE III An alloy of vanadium, comprising 7.48 percent aluminum and 0.44 percent oxygen, was prepared by the method described in Example I except that the mix ratio was varied from stoichiometric. Approximately 908 grams vanadium pentoxide and 500 grams of aluminum were reacted to produce 465 grams of this alloy at a yield of 84.5 per cent. The alloy was crushed to Ai-inch mesh and through prior to feeding 217 grams of it into a fused electrolyte of sodium chloride and subvalent vanadium chloride, likewise prepared as described in Example I. Analysis showed the electrolyte to contain 6.2 percent soluble vanadium. A steel cathode was immersed into the electrolyte to a depth of one inch, and amperes of direct current were passed for 5 hours and 55 minutes. The deposit contained 21.024 grams of leached metal indicating a current efliciency of 90 percent based on V. The metal analyzed: 0.28 percent oxygen and 0.68 percent aluminum. A second electrolysis with no further addition was run at a current of 5 amperes for 4 hours and 30 minutes. After leaching the vanadium, the deposit weighed 18.09 grams, indicating a current efficiency of 83.3 percent. This metal was arc-melted into button form which had a Rockwell A hardness of 40.6. Chemical analysis showed 0.10 percent oxygen and 0.048 percent aluminum. A third electrolysis was run at 5 amperes for 4 hours and 50 minutes to give 19.9 grams of metal at a current efficiency of 88.4 percent which analyzed: 0.01 percent oxygen, 0.09 percent aluminum, and had a Rockwell A hardness of 45. Before starting the fourth electrolysis period, 136 grams of electroltye analyzing 6.2 percent soluble vanadium were added to the cell to increase the depth of the bath. Electrolysis was continued with a current of 5 amperes at 0.37 volt for 3 hours and 50 minutes, 15.4 grams of metal deposited at a current efliciency of 83.4 percent. This metal analyzed: 0.1 percent oxygen and 0.21 percent aluminum. For the fifth electrolytic period, current was passed at a rate of 5 amperes at 0.32 volt for a period of 4 hours and 40 minutes, depositing on the cathode 20.6 grams of metal at a current efliciency of 95 percent based on divalent vanadium. This metal analyzed: 0.07 percent oxygen and 0.10 percent aluminum. This electrolysis was continued for 6 additional electrolytic periods in which 5 amperes of current were passed for a total of 26 hours and 12 minutes. Eighty-seven and six tenths grams of metal were deposited at an average current efiiciency of 68.7. A sum total of 181.9 grams of vanadium metal were deposited on the cathode out of the 200 grams available in the 217 grams of anode source material during the run. This amounted to approximately 91 percent vanadium recovery. The residue that was leached from the electrolyte analyzed: 20.71 percent aluminum, 7.8 percent oxygen, and 0.69 percent nitrogen. X-ray diffraction pattern indicated the residue (anode sludge) to contain A1 0 Spectrographic analysis of the residues (anode sludge) indicated that it was composed of chiefly vanadium and aluminum. Large quantities of nickel, silicon, and small quantities of molybdenum were also detected.

EXAMPLE IV Two cells connected in series were each charged with 200 grams of vanadium-aluminum alloy (analyzing 3.8 percent aluminum, 0.39 percent oxygen) and 850 grams of NaCl-VC1 electrolyte (7 percent VCl Five grams of aluminum shot were added to one of these cells. The cells were then prepared as previously described and electrolyzed for 25 ampere-hours. At the completion of electrolysis and cooling, the two cathodes were compared. The one to which the live grams of free aluminum has been added exhibited large crystals. The cathode deposit from the cell to which no aluminum had been added did not have the large crystal growth. The obvious improvement brought about by the addition of the free aluminum was Verified by chemical analysis as shown in the Table.

Table I FREE ALUMINUM ADDITIONS This test showed that the excess aluminum, added independently as a scavenger or as a constituent of the anodic alloy, competes successfully with the cathode deposit for the oxygen and nitrogen in the system. However, the addition of too much aluminum to the electrolyte could reduce its vanadium concentration below that necessary for satisfactory electrorefining.

From the above description, it will be seen that a novel process has been disclosed by which a vanadium metal, free from embr-ittling elements, can be produced more economically than heretofore possible with prior art methods.

What is claimed is:

1. In a process for the electrorefining of oxygen-contaminated vanadium in an electrolytic cell wherein an electrode and the vanadium to be refined are immersed in, and separated from each other by, an electrolyte comprising a lower vanadium halide dissolved in fused electrolytic salt, and said electrode is rendered cathodic and said vanadium is rendered anodic under the influence of direct current applied to said electrolytic cell to efiect the deposition of vanadium from said electrolyte on said electrode and the dissolution of the anodic vanadium in said electrolyte, the improvement for producing vanadium with substantially reduced oxygen content which comprises introducing elemental aluminum into said electrolyte in electrical relation with said vanadium to render said aluminum anodic and electrodepositing vanadium from said electrolyte on the cathodic electrode.

2. A process in accordance with claim 1 wherein said aluminum is introduced into said electrolyte in an amount such that the ratio of aluminum to the oxygen contamination of the vanadium to be refined is at least that of the ratio of aluminum to oxygen in aluminum sesquioxide and less than 10 percent in excess of that ratio.

3. A process for the production of ductile vanadium having a low oxygen contamination which comprises mixing oxidic pentavalent vanadium values with at least the stoichiometric amount of aluminum for the aluminum sesquioxide-and-elemental vanadium-forming reaction and less than 10 percent in excess of the stoichiometric requirement, heating said mixture to the ignition temperature for the reduction of said oxidic vanadium by said aluminum to produce an alloy containing vanadium, a minor proportion of aluminum, balance incidental impuri- 7 ties, introducing said vanadium-aluminum alloy into an electrorefining cell having a cathode immersed in an electrolyte comprising a fused electrolytic salt having dissolved therein a lower halide of vanadium, passing direct current through said electrorefining cell in a manner to render said vanadium alloy anodic with respect to said cathode whereby the vanadium in said vanadium alloy is dissolved in said electrolyte and electrodepositing the dissolved vanadium on said cathode in a substantially oxygen-free form, and separating the cathode product from 10 the electrolyte.

4. A process in accordance with claim 3 wherein the 8 vanadium alloy produced in the aluminum reduction step contains vanadium, aluminum in an amount of at least 2 percent and less than 7.5 percent, balance incidental impurities.

References Cited in the file of this patent UNITED STATES PATENTS 2,789,896 Cofier Apr. 23, 1957 2,913,380 Gullett NOV. 17, 1959 FOREIGN PATENTS 817,893 Great Britain Aug. 6, 1959 

1. IN A PROCESS FOR THE ELECTROREFINING OF OXYGEN-CONTAMINATED VANADIUM IN AN ELECTROLYTIC CELL WHEREIN AN ELECTRODE AND THE VANADIUM TO BE REFINED ARE IMMERSED IN, AND SEPARATED FROM EACH OTHER BY, AN ELECTROLYTE COMPRISING A LOWER VANADIUM HALIDE DISSOLVED IN FUSED ELECTROLYTIC SALT, AND SAID ELECTRODE IS RENDERED CATHODIC AND SAID VANADIUM IS RENDERED ANODIC UNDER THE INFLUENCE OF DIRECT CURRENT APPLIED TO SAID ELECTROLYTIC CELL TO EFFECT THE DEPOSITION OF VANADIUM FROM SAID ELECTROLYTE ON SAID ELECTRODE AND THE DISSOLUTION OF THE ANODIC VANADIUM IN SAID ELECTROLYTE, THE IMPROVEMENT FOR PRODUCING VANADIUM WITH SUBSTANTIALLY REDUCED OXYGEN CONTENT WHICH COMPRISES INTRODUCING ELEMENTAL ALUMINUM INTO SAID ELECTROLYTE IN ELECTRICAL RELATION WITH SAID VANADIUM TO RENDER SAID ALUMINUM ANODIC AND ELECTRODESPOSITING VANADIUM FROM SAID ELECTROLYTE ON THE CATHODIC ELECTRODE. 